Method for thiosulfate leaching of precious metal-containing materials

ABSTRACT

Processes are provided for recovering precious metals from refractory materials using thiosulfate lixiviants. The processes can employ lixiviants that include at most only small amounts of copper and/or ammonia and operate at a relatively low pH, reduction of polythionates, inert atmospheres to control polythionate production, and electrolytic solutions which provide relatively high rates of precious metal recovery.

CROSS REFERENCE TO RELATED APPLICATION

[0001] The present application claims priority from U.S. ProvisionalApplication Ser. No. 60/205,472, filed May 19, 2000, which isincorporated herein by this reference.

FIELD OF THE INVENTION

[0002] The present invention is directed generally to the recovery ofprecious metals from precious metal-containing material and specificallyto the recovery of precious metals from precious metal-containingmaterial using thiosulfate lixiviants.

BACKGROUND OF THE INVENTION

[0003] A traditional technique for recovering precious metal(s) fromprecious metal-containing ore is by leaching the material with a cyanidelixiviant. As used herein, a “precious metal” refers to gold, silver,and the platinum group metals (e.g., platinum, palladium, ruthenium,rhodium, osmium, and iridium). Many countries are placing severelimitations on the use of cyanide due to the deleterious effects ofcyanide on the environment. Incidents of fish and other wildlife havingbeen killed by the leakage of cyanide into waterways have been reported.The limitations being placed on cyanide use have increased substantiallythe cost of extracting precious metal(s) from ore, thereby decreasingprecious metal reserves in many countries. Cyanide is also unable torecover precious metals such as gold from refractory ores without apretreatment step. “Refractory ores” refer to those ores that do notrespond well to conventional cyanide leaching. Examples of refractoryores include sulfidic ores (where at least some of the precious metalsare locked up in the sulfide matrix), carbonaceous ores (where theprecious metal complex dissolved in the lixiviant adsorbs ontocarbonaceous matter in the ores), and sulfidic and carbonaceous ores.

[0004] Thiosulfate has been actively considered as a replacement forcyanide. Thiosulfate is relatively inexpensive and is far less harmfulto the environment than cyanide. Thiosulfate has also been shown to beeffective in recovering precious metals from pretreated refractorypreg-robbing carbonaceous ores and sulfidic ores. As used herein,“preg-robbing” is any material that interacts with (e.g., adsorbs orbinds) precious metals after dissolution by a lixiviant, therebyinterfering with precious metal extraction, and “carbonaceous material”is any material that includes one or more carbon-containing compounds,such as humic acid, graphite, bitumins and asphaltic compounds.

[0005] Where gold is the precious metal, thiosulfate leaching techniqueshave typically relied on the use of copper ions to catalyze andaccelerate the oxidation of gold, ammonia to facilitate the formationand stabilization of cupric ammine ions and/or a pH at pH 9 or above tomaintain a region of stability where both the cupric ammine and goldthiosulfate complexes are stable.

[0006] It is well known in the art that the catalytic effect of copperand ammonia in conventional thiosulfate leaching of gold is described bythe following sequence of reactions. Formation of the cupric amminecomplex:

Cu ²⁺ +4NH³→Cu(NH³)⁴ ²⁺   (1)

[0007] Oxidation of gold by cupric ammine, gold complexation as thegold-thiosulfate anion, and reduction of the cupric ammine to cuprousthiosulfate:

Au+Cu(NH₃)₄ ²⁺+5S₂O₃ ²⁻→Au(S₂O₃)₂ ³⁻+Cu(S₂O₃)₃ ⁵⁻+4NH₃  (2)

[0008] Oxidation of the cuprous thiosulfate back to cupric ammine withoxygen:

Cu(S₂O₃)₃ ⁵⁻+4NH₃+¼O₂+½H₂O→Cu(NH₃)₄ ²⁺+3S₂O₃ ²⁻+OH⁻  (3)

[0009] Summing equations (2) and (3) yields the overall thiosulfateleach reaction for gold:

Au+2S₂O₃ ²⁻+¼O₂+H₂O→Au(S₂O₃)₂ ³⁻+OH⁻  (4)

[0010] It can be seen from the above equations that copper and ammoniaact as catalysts in that they are neither produced nor consumed in theoverall leach reaction.

[0011] Cupper and ammonia can be a source of problems. Added coppertends to precipitate as cupric sulfide, which is speculated to form apassive layer on gold, thereby inhibiting gold leaching as well asincreasing copper and thiosulfate consumption:

Cu²+S₂O₃ ²⁻+2OH⁻→CuS+SO₄ ²⁻+H₂O  (5)

[0012] Rapid oxidation of thiosulfate by cupric ammine also occurs,leading to excessive degradation and loss of thiosulfate:

2Cu(NH₃)₄ ²⁺+8S₂O₃ ²⁻→2Cu(S₂O₃)₃ ⁵⁻+S₄O₆ ²⁻+8NH₃  (6)

[0013] Loss of ammonia by volatilization occurs readily, particularly inunsealed gas-sparged reactors operating at pH greater than 9.2, leadingto excessive ammonia consumption:

NH₄ ⁺+OH⁻→NH_(3(aq))+H₂O→NH_(3(g))+H₂O  (7)

[0014] Like cyanide, copper and ammonia are highly toxic to many aquaticlifeforms and are environmentally controlled substances.

[0015] Other problems encountered with thiosulfate leaching includedifficulty in recovering gold out of solution as a result of theformation of polythionates, such as tetrathionate and trithionate, whichadsorb competitively with gold onto adsorbents, such as resins. Theformation of polythionates further increases thiosulfate consumption perunit mass of processed ore.

SUMMARY OF THE INVENTION

[0016] These and other needs have been addressed by the methodologiesand systems of the present invention. The methodologies can recoverprecious metals from a variety of materials, including refractorycarbonaceous or sulfidic ores, double refractory ores (e.g., orescontaining both sulfide-locked gold and carbonaceous preg-robbingmatter), oxide ores, nonrefractory sulfidic ores, and ores alsocontaining copper minerals and other materials derived from such ores(e.g., concentrates, tailings, etc.).

[0017] In one embodiment, a thiosulfate leaching process is providedthat includes one or more of the following operating parameters:

[0018] (a) an oxygen partial pressure that is preferablysuperatmospheric and more preferably ranges from about 4 to about 500psia;

[0019] (b) a leach slurry pH that is preferably less than pH 9;

[0020] (c) a leach slurry that is preferably at least substantially freeof (added) ammonia and more preferably contains less than 0.05M (added)ammonia such that the leach slurry has a maximum total concentration ofammonia of preferably less than 0.05M and more preferably no more thanabout 0.025M;

[0021] (d) a leach slurry that is preferably at least substantially freeof (added) copper ion and more preferably contains no more than about 15ppm (added) copper ions;

[0022] (e) an (added) sulfite concentration that is preferably no morethan about 0.01 M such that the slurry has a maximum total concentrationof sulfite of preferably no more than about 0.02M and more preferably nomore than about 0.01M; and/or (f) a leach slurry temperature preferablyranging from about 20 to about 100° C. and more preferably from about 20to about 80° C. The foregoing parameters can yield a high level ofprecious metal extraction from the precious metal-containing material,which can be at least about 70% and sometimes at least about 80%.

[0023] The thiosulfate lixiviant can be derived from any suitableform(s) ofthiosulfate, such as sodium thiosulfate, calcium thiosulfate,potassium thiosulfate and/or ammonium thiosulfate. Sodium and/or calciumthiosulfate are preferred.

[0024] The leaching process can be conducted by any suitable technique.For example, the leaching can be conducted in situ, in a heap or in anopen or sealed vessel. It is particularly preferred that the leaching beconducted in an agitated, multi-compartment reactor such as anautoclave.

[0025] The precious metal can be recovered from the pregnant leachsolution by any suitable technique. By way of example, the preciousmetal can be recovered by resin adsorbtion methods such asresin-in-pulp, resin-in-solution, and resin-in-leach or by solventextraction, cementation, electrolysis, precipitation, and/orcombinations of two or more of these techniques.

[0026] Reducing or eliminating the need to have copper ions and/orammonia present in the leach as practiced in the present invention canprovide significant multiple benefits. First, the cost of having to addcopper and ammonia reagents to the process can be reduced significantlyor eliminated. Second, environmental concerns relating to the presenceof potentially harmful amounts of copper and ammonia in the tailings orother waste streams generated by the process can be mitigated. Third,the near-absence or complete absence of copper and ammonia in the leachcan provide for a much more reliable and robust leaching process,yielding more stable leachates, able to operate over a wider pH andoxidation-reduction potential (ORP) range than is possible withconventional thiosulfate leaching. The latter process must operate inthe relatively narrow window of pH and ORP where both the cupric amminecomplex and the gold thiosulfate complex co-exist. With the process ofthe present invention, the pH of the thiosulfate lixiviant solution inthe leaching step can be less than pH 9 and the ORP less than 200 mV(referenced to the standard hydrogen electrode). Fourth, minimizing theamount of copper in the system can lead to increased loading of goldonto resins due to reduced competitive adsorption of copper ions. Resinelutions are also simplified as little, if any copper, is on the resin.Finally, the near-absence or complete absence of copper and ammonia inthe leach can reduce or eliminate entirely a host of deleterious sidereactions that consume thiosulfate and are otherwise difficult orimpossible to prevent.

[0027] The elimination or near elimination of sulfite from thethiosulfate leach also can have advantages. Sulfite can depress the rateof dissolution of precious metal from the precious metal-containingmaterial by reducing significantly the oxidation reduction potential(ORP) of the leach solution or lixiviant. As will be appreciated, therate of oxidation of the gold (and therefore the rate of dissolution ofthe gold) is directly dependent on the ORP.

[0028] In another embodiment, an extraction agent is preferablycontacted with a pregnant (precious metal-containing) thiosulfate leachsolution at a temperature of less than about 70° C. and more preferablyless than about 60° C. in the substantial absence of dissolved molecularoxygen to isolate the precious metal and convert polythionates in thepregnant leach solution into thiosulfate. In one configuration, theextraction agent is an adsorbent, such as a resin, which loads theprecious metal onto the adsorbent. As used herein, an “adsorbent” is asubstance which has the ability to hold molecules or atoms of othersubstances on its surface. Examples of suitable resin adsorbents includeweak and strong base resins such as “DOWEX 21K”, manufactured by DowChemical. In another configuration, the extraction agent is a solventextraction reagent that extracts the precious metals into an organicphase, from which the precious metals can be later recovered. As will beappreciated, the detrimental polythionates decompose into thiosulfate inthe substantial absence of dissolved molecular oxygen.

[0029] In yet another embodiment, the pregnant leach solution from athiosulfate leaching step is contacted, after the leaching step, with areagent to convert at least about 50% and typically at least most ofpolythionates (particularly trithionate and tetrathionate) intothiosulfate. The reagent or reductant can be any suitable reactant toconvert polythionates into thiosulfate, with any sulfide, and/orpolysulfide (i.e., a compound containing one or a mixture of polymericion(s) S_(x) ²⁻, where x=2-6, such as disulfide, trisulfide,tetrasulfide, pentasulfide and hexasulfide) being particularlypreferred. A sulfite reagent can also be used but is generally effectiveonly in converting polythionates of the form S_(x)O₆ ²⁻, where x=4 to 6,to thiosulfate. The sulfite, sulfide, and/or polysulfide can becompounded with any cation, with Groups IA and IIA elements of thePeriodic Table, ammonium, and hydrogen being preferred.

[0030] In yet another embodiment, a precious metal solubilized in asolution, such as a pregnant leach solution or eluate, is electrowon inthe presence of sulfite. In the presence of sulfite, the precious metalis reduced to the elemental state at the cathode while the sulfite isoxidized to sulfate at the anode. Sulfite is also believed to improvethe precious metal loading capacity of the resin by converting loadedtetrathionate to trithionate and thiosulfate.

[0031] In yet another embodiment, the formation of polythionates iscontrolled by maintaining a (pregnant or barren) thiosulfate leachsolution in a nonoxidizing (or at least substantially nonoxidizing)atmosphere and/or sparging a nonoxidizing (or at least substantiallynonoxidizing) gas through the leach solution. As will be appreciated,the atmosphere or gas may contain one or more reductants, such ashydrogen sulfide and/or sulfur dioxide. The molecular oxygenconcentration in the atmosphere and/or sparge gas is preferablyinsufficient to cause a dissolved molecular oxygen concentration in theleach solution of more than about 1 ppm and preferably of more thanabout 0.2 ppm. Preferably, the inert atmosphere (or sparge gas) is atleast substantially free of molecular oxygen and includes at least about85 vol. % of any inert gas such as molecular nitrogen and/or argon. Bycontrolling the amount of oxidant(s) (other than thiosulfate andpolythionates) in the atmosphere and/or (pregnant or barren) leachsolution the rate or degree of oxidation of thiosulfates to formpolythionates can be controlled.

BRIEF DESCRIPTION OF THE DRAWINGS

[0032]FIG. 1 is a flow schematic of a first embodiment of the presentinvention;

[0033]FIG. 2 is a flow schematic of second embodiment of the presentinvention;

[0034]FIG. 3 is a flow schematic of a third embodiment of the presentinvention;

[0035]FIG. 4 is a flow schematic of a fourth embodiment of the presentinvention;

[0036]FIG. 5 is a plot of gold extraction in percent (vertical axis)versus leach time in hours (horizontal axis);

[0037]FIG. 6 is another plot of gold extraction in percent (verticalaxis) versus leach time in hours (horizontal axis);

[0038]FIG. 7 is another plot of gold extraction in percent (verticalaxis) versus leach time in hours (horizontal axis);

[0039]FIG. 8 is another plot of gold extraction in percent (verticalaxis) versus leach time in hours (horizontal axis); and

[0040]FIG. 9 is a plot of gold extraction in percent (left verticalaxis) and thiosulfate remaining in percent (right vertical axis) versusleach time in hours (horizontal axis).

DETAILED DESCRIPTION

[0041] The present invention provides an improved thiosulfate leachingprocess for the recovery of precious metals from precious metal-bearingmaterial. The precious metal(s) can be associated with nonpreciousmetals, such as base metals, e.g., copper, nickel, and cobalt. Theprecious metal-bearing material includes ore, concentrates, tailings,recycled industrial matter, spoil, or waste and mixtures thereof. Theinvention is particularly effective for recovering precious metals,particularly gold, from refractory carbonaceous material.

[0042]FIG. 1 is a flow chart according to a first embodiment of thepresent invention. The process of the flow chart is particularlyeffective in recovering gold from carbonaceous material and oxidematerial and mixtures thereof.

[0043] Referring to FIG. 1, aprecious metal-bearing material 100 issubjected to the steps of wet and/or dry crushing 104 and wet and/or drygrinding 108 to reduce the particle size of the material sufficiently toenable the solids to be suspended in an agitated vessel and to allow forthe efficient leaching of the precious metals. Preferably, wet grindingis employed with the recycled thiosulfate leach solution and water beingused as the liquid component in the slurry. In that event, the slurry112 containing the comminuted material typically contains from about0.05 to about 0.1 M thiosulfates and from about 0.0005 to about 0.025 Mpolythionates. The fully comminuted material particle size is preferablyat least smaller than 80% passing about 48 mesh (300 microns), morepreferably 80% passing about 100 mesh (150 microns), and most preferably80% passing about 200 mesh (75 microns). The typical solids content ofthe slurry 112 ranges from about 20 to about 30 wt. %. As will beappreciated, other techniques can be used to comminute the material tothe desired particle size(s). By way of illustration, blasting can beused alone with or without crushing and grinding and crushing andgrinding can be used alone with or without another comminutiontechnique.

[0044] The ground slurry 112 is then thickened 116 to adjust the pulpdensity to a value suitable for leaching. The ideal leach pulp densitywill vary according to the type of material being leached. Typically,the pulp density ranges from about 20 to about 50% solids by weight, butcould be as low as about 1% or as high as about 60%. Thickening 116 willgenerally not be required if the desired pulp density (after wetcomminution or formation of the comminuted material into a slurry) isless than about 20%.

[0045] The thickener overflow solution 120 is recycled back to grinding108 in the event that wet grinding is employed. Otherwise, the overflowsolution 120 is returned to the optional slurry formation step (notshown).

[0046] Fresh makeup thiosulfate is added, as necessary, at any suitablelocation(s), such as to the slurried material during comminution 108and/or in the thickener 116, to the underflow or overflow solution 124,120, to leaching 132 and/or to the regenerated thiosulfate solution 128(discussed below). In any event, the optimum solution thiosulfateconcentration to maintain during leaching 132 will depend on the natureof the material being leached, but will preferably range from about0.005 to about 2 molar (M), more preferably about 0.02 to about 0.5 M,and even more preferably from about 0.05 to about 0.2 M. The source ofmakeup thiosulfate can be any available thiosulfate-containing compound,such as sodium thiosulfate, potassium thiosulfate, calcium thiosulfate,or any other thiosulfate-containing material or thiosulfate precursor.Ammonium thiosulfate can also be used but its use is less preferred forenvironmental reasons. Alternatively, thiosulfate can be generated insitu or in a separate step by reaction of elemental sulfur with a sourceof hydroxyl ions, in accordance with the following reaction:

2(x+1)S+60H⁻→S₂O₃ ²⁻+2S_(x) ²⁻+3H₂O  (8)

[0047] where x=3-6, or by reaction of bisulfide with bisulfite:

2HS⁻+4HSO₃ ⁻→3S₂O₃ ²⁻+3H₂O  (9)

[0048] or by reaction of elemental sulfur with sulfite:

S+SO₃ ²⁻→S₂O₃ ²⁻  (10)

[0049] If the desirable temperature is above ambient, it may bedesirable to recover waste heat for recycle to leaching. In that event,the underflow slurry 124 is directed through an indirect heat exchanger136, preferably a shell and tube heat exchanger system in which the hotslurry from resin-in-pulp pretreatment 140 (discussed below) is passedthrough the inner tubes and the cold feed (or underflow) slurry 140 ispassed through the annular space between the tubes (or vice versa). Inthis way waste heat is transferred from the leached slurry 144 to thefeed (or underflow) slurry 124, reducing the amount of new heat thatmust be added in leaching 132 to maintain the desired leach temperature.Typically, the approach temperature of the incoming feed slurry 148 isfrom about 2 to about 5° C. below the leach temperature (discussedbelow) and heat is added to the leach vessel by suitable techniques tomakeup the difference.

[0050] The heated slurry 148 is subjected to leaching 132 in thepresence of oxygen and thiosulfate. Leaching is conducted in thepresence of an oxygen-enriched atmosphere at atmospheric pressure, or ata pressure above atmospheric pressure using an oxygen-containing gas toreduce or eliminate the need for the presence of copper and/or ammoniain the leach. Using gold as an example, the thiosulfate leaching ofprecious metal-bearing material in the absence or substantial absence ofcopper and ammonia under elevated oxygen partial pressure can beillustrated by the following reaction:

Au+2S₂O₃ ²⁻+¼O₂+½H₂O→Au(S₂O₃)₂ ³⁻+OH⁻  (11)

[0051] The increased oxygen partial pressure in the leach increases therate of the above reaction in the absence or near absence of copper andammonia. To accomplish this goal, the oxygen-containing gas may includeatmospheric air, or it may include relatively pure (95%+) oxygen such asthat produced from any commercially available oxygen plant, or it mayinclude any other available source of oxygen. The desired oxygen partialpressure (PO₂) maintained during leaching will depend on the materialbeing leached, but it will be at least higher than that provided undernormal ambient conditions by air at the elevation the process isapplied. Thus, if the process is practiced at sea level for example theoxygen partial pressure will be in excess of about 3 pounds per squareinch absolute pressure (psia) to as high as about 500 psia, preferablyfrom about 10 to about 115 psia, and most preferably from about 15 toabout 65 psia. The total operating pressure is the sum of the molecularoxygen partial pressure and the water vapor pressure at the temperatureemployed in the leaching step 132, or preferably ranges from about 15 toabout 600 psia and more preferably from about 15 to about 130 psia.

[0052] The leaching temperature will be dictated by the type of materialbeing leached. The temperature will vary typically from about 5° C. toabout 150° C., preferably from about 20 to about 100° C., and mostpreferably from about 40 to about 80° C. Higher temperatures acceleratethe leaching of precious metals but also accelerate the degradation ofthiosulfate. If required, a source of makeup heat such as steam is addedto the leach reactors to maintain the desired temperature.

[0053] The leaching retention time is dependent on the material beingleached and the temperature, and will range from about 1 hour to 96hours, preferably from about 2 to about 16 hours, and most preferablyfrom about 4 to about 8 hours.

[0054] The absence or substantial absence of copper and/or ammonia inthe leach greatly simplifies the process. Elimination ornear-elimination of ammonia and copper from the leach provides theadvantage of allowing for a consistently high and reproducible preciousmetal extraction over a broader pH range than was previously possiblewith the other thiosulfate leaching processes. Preferably, the (addedand/or total solution) copper concentration is no more than about 20ppm, more preferably no more than about 15 ppm, and even more preferablyno more than about 10 ppm while the (added and/or total solution)ammonia concentration is no more than about 0.05 M, more preferably nomore than about 0.03 M, and even more preferably no more than about 0.01M. In the present invention leaching can be operated at about pH 7-12,preferably about pH 8-11, more preferably about pH 8-10, and even morepreferably at a pH less than pH 9. The oxidation-reduction potential(ORP) preferably ranges from about 100 to about 350 mV and morepreferably from about 150 to about 300 mV (vs. the standard hydrogenelectrode (SHE)).

[0055] Oxidative degradation of thiosulfate ultimately to sulfate canalso occur, possibly by the following sequence of reactions that involvethe formation of intermediate polythionates (polythionates can berepresented by S_(n)O₆ ²⁻, where n=2-6):

Tetrathionate formation: 2S₂O₃ ²⁻+½O₂+H₂O→S₄O₆ ²⁻+2OH  (12)

Trithionate formation: 3S₄O₆ ²⁻ +{fraction (5/2)}O ₂+H₂O→4S₃O₆²⁻+2H⁺  (13)

Sulfite formation: S₃O₆ ²⁻+½O₂+2H₂O→3SO₃ ²⁻+4H⁺  (14)

Sulfate formation: 2SO₃ ²⁻+O₂→2SO₄ ²⁻  (15)

Overall: S₂O₃ ²⁻+2O₂+H₂O→2SO₄ ²⁻+2H⁺  (16)

[0056] Oxidative degradation of thiosulfate to polythionates andsulfates is accelerated markedly in the presence of copper ions and/orammonia. The oxidative degradation reactions are slowed considerably atelevated oxygen partial pressure in the absence or near-absence ofcopper and ammonia.

[0057] The leaching step 132 may be conducted in a batch or continuousbasis but continuous operation is preferred. Continuous leaching iscarried out in a multiple series of one or more reactors that areagitated sufficiently to maintain the solids in suspension. Agitationmay be accomplished by mechanical, pneumatic or other means. In apreferred configuration, gassing impellers are employed, such as thosedisclosed in U.S. Pat. No. 6,183,706 and copending U.S. patentapplication Ser. No. 09/561,256, filed Apr. 27, 2000, which areincorporated herein by reference. Such impellers can significantlyenhance the amount of dissolved molecular oxygen in the leach solution.Leaching may also be carried out in a multi-compartment autoclavecontaining one or more compartments, (with 4 to 6 compartments beingpreferred) similar in design to the autoclaves used to pressure oxidizesulfide-bearing ores or concentrates. However, owing to the non-acidic,moderate temperature, relatively mild conditions employed in the presentinvention, the autoclave materials of construction are much lessexpensive than those found to be necessary when oxidizing sulfideminerals. The latter autoclaves are normally constructed of a steelshell fitted with a lead liner and refractory brick liner and containingmetallic components constructed of titanium or other expensivecorrosion-resistant alloys. The leach reactors and contained metalliccomponents employed by the present invention can be simply constructedof stainless steel and do not require lead or brick liners.

[0058] The extraction of precious metals in the leaching step 132 isrelatively high, particularly for carbonaceous ores. Typically, at leastabout 50%, more typically at least about 70%, and even more typically atleast about 80% of the precious metal in the precious metal-containingmaterial is extracted or solubilized into the pregnant leach solution144. The concentration of the dissolved precious metal in the pregnantleach solution typically ranges from about 0.05 to about 100 ppm andmore typically from about 1 to about 50 ppm.

[0059] The pregnant leach slurry 144 containing the preciousmetal-bearing leach solution and gold-depleted solid residue mayoptionally be directed to RIP pretreatment 140 to reduce theconcentration of polythionates in solution. As will be appreciated, themolecular oxygen sparged through the leach slurry in the leaching step132 will oxidize a minor portion of the thiosulfate into polythionates.Polythionates have the undesired effect of reducing the loading ofprecious metals on to resin by competitive adsorption. Lowering thepolythionate concentration will have the beneficial effect of increasingthe loading of precious metals on to resin, thereby improving theefficiency of resin recovery of precious metals.

[0060] The RIP pretreatment step 140 can be performed using any one ormore of a number of techniques for converting polythionates to othercompounds that do not compete with the precious metal for collection bythe extraction agent.

[0061] In one embodiment, a polythionate reductant is added to theslurry 144 to reduce polythionates to thiosulfates. Any of a number ofreductants are suitable for performing the conversion.

[0062] By way of example, a sulfide-containing reagent can reduce thepolythionates back to thiosulfate, as shown by the following reactions:

2S₄O₆ ²⁻+S²⁻+{fraction (3/2)}H₂O→{fraction (9/2)}S₂O₃ ²⁻+3H⁺  (17)

S₃O₆ ²⁻+S²⁻→2S₂O₃ ²⁻  (18)

[0063] Any reagent that releases sulfide ions on dissolution willsuffice, such as sodium bisulfide, NaHS, sodium sulfide, Na₂S, hydrogensulfide gas, H₂S, or a polysulfide. The use of a sulfide reagent has theadvantages of rapidly and efficiently reducing polythionates tothiosulfate at ambient or moderately elevated temperature. The treatmentcan be carried out in an agitated reactor in batch mode or in a seriesof 1-4 reactors operating in continuous mode, or in a multi-compartmentautoclave. Alternatively the treatment can be carried out in a pipereactor or simply by injecting sulfide ions in the piping systemdirecting the leach slurry to gold recovery, or the first stage of RIP.The treatment is carried out at a controlled pH of about pH 5.5 to aboutpH 10.5, preferably about pH 7 to about pH 10, most preferably less thanabout pH 9. The temperature employed can range from about 20° C. toabout 150° C., preferably from about 40 to about 100° C., morepreferably from about 40 to about 80° C., and even more preferably fromabout 60 to about 80° C. The retention time can range from as low as 5minutes, preferably greater than 30 minutes, most preferably from about1 to about 3 hours.

[0064] Alternatively, a sulfite-containing reagent can also reducepolythionates to thiosulfates as shown by the following reaction:

S₄O₆ ²⁻+SO₃ ²⁻→S₂O₃ ²⁻+S₃O₆ ²⁻  (19)

[0065] Sulfite treatment is effective in reducing tetrathionate quickly,but a disadvantage is it is ineffective in reducing trithionate. Thesulfite can be added in any form and/or can be formed in situ. Forexample, sulfite can be added in the form of sodium metabisulfite orsulfur dioxide.

[0066] When using sulfite, the temperature of the leach slurry in theRIP pretreatment 140 is preferably less than 60° C. to inhibit, at leastsubstantially, the precipitation of precious metal(s) from the leachslurry 144. More preferably, the RIP pretreatment 140 with sulfite isperformed at a temperature in the range of from about 10 to about 50° C.and even more preferably at ambient temperature.

[0067] When using sulfite, the residence time of the leach slurry 144 inthe regeneration step 140 is preferably at least about 1 second, morepreferably greater than about 5 minutes, and even more preferablygreater than about 10 minutes and no more than about 1 hour, with about15-30 minutes being most preferable.

[0068] The pH of the leach slurry during sulfite treatment typicallyranges from about pH 5.5 to about pH 10.5 and more typically from aboutpH 7 to about pH 10.

[0069] Other suitable polythionate reductants include hydrogen, fine,reactive elemental sulfur, carbon monoxide, and mixtures thereof.

[0070] In another embodiment, the pretreatment step 140 is performed bymaintaining the temperature of the leach slurry at a sufficiently highvalue in the absence of oxygen to effect the following hydrolyticdisproportionation reactions:

4S₄O₆ ²⁻+5H₂O→7S₂O₃ ²⁻+2SO₄ ²⁻+10H⁺  (20)

S₃O₆ ²⁻+H₂O→S₂O₃ ²⁻+SO₄ ²⁻+2H⁺  (21)

[0071] Hydrolytic treatment can be carried out in an agitated reactor inbatch mode or in a series of 1-4 reactors operating in continuous mode,or a multi-compartment autoclave. The temperature is preferablymaintained in the range of from about 60 to about 150° C., preferably offrom about 70 to about 100° C., and most preferably of from about 80 toabout 90° C. by adding a source of heat, such as steam. The retentiontime typically ranges from about 15 minutes to 8 hours, preferably fromabout 1 to about 6 hours, and most preferably from about 2 to about 4hours. Hydrolytic treatment is generally less preferable than sulfidetreatment because the former method results in irreversible loss of someof the polythionate to sulfate.

[0072] Alternatively, any or all of the above-techniques for convertingpolythionate(s) into thiosulfate can be combined in the same processconfiguration.

[0073] In a preferred embodiment, the reductant used to convertpolythionates into thiosulfates is the sulfide reagent. Sulfide additionis preferred because one sulfide reacts with one tri- or twotetrathionates to form multiple thiosulfates without anysulfur-containing byproducts. Sulfite addition only reducestetrathionate and is not capable of reducing trithionate at commonoperating temperatures and pH's. Heating of the leach solution is energyintensive and produces byproducts. Trithionate and tetrathionate areeach converted into thiosulfate, sulfate, and hydrogen ions, thus thethiosulfate yield is not as high as with sulfide addition.

[0074] RIP pretreatment 140 can be performed in any suitable vessel(s),preferably agitated. Preferably, RIP pretreatment is performed in aseries of tanks or in a multistaged vessel.

[0075] The addition of a sulfide such as NaHS is preferred. Preferably,the amount of the reductant generally and, sulfide reagent specificallyadded to the slurry 144 is sufficient to convert at least most of thepolythionates into thiosulfate. The amount of sulfide contacted with theslurry 144 preferably is at least about 100 to about 150% of thestoichiometric amount required to convert at least substantially all ofthe polythionates in the slurry into thiosulfates. Typically, at leastabout 50%, more typically at least most, and even more typically fromabout 80 to about 95% of the polythionates are converted intothiosulfates in RIP pretreatment 140.

[0076] The temperature of the slurry 144 preferably is at least about60° C. and the ORP of the exiting slurry 152 is at least below about 100mV (SHE) and more preferably ranges from about −100 to about 100 mV(SHE) to substantially minimize precious metal precipitation.

[0077] The exiting RIP pretreated slurry 152 is passed through heatexchanger 136 and conditioned in a conditioner 156 to resolubilize anyprecious metal precipitated during RIP pretreatment 140 and/or heatexchange 136. Conditioning 156 is performed in an agitated single- ormulti-compartment vessel which has an oxidizing atmosphere, such as air,to cause solubilization of the precious metal precipitates. Althoughpolythionates will form in the presence of an oxidant, such as molecularoxygen, the rate of conversion of thiosulfate to polythionates is muchslower than the rate of precious metal solubilization. Preferably, theresidence time (at ambient temperature and pressure) is selected suchthat at least about 95 % of the precious metal precipitates aresolubilized while no more than about 5% of the thiosulfate is convertedinto polythionates. Preferably, the slurry residence time inconditioning 156 is no more than about 12 hrs and more preferably rangesfrom about 1 to about 6 hrs.

[0078] The conditioned slurry 160 is next subjected to resin-in-pulptreatment 164 to extract the precious metal from the conditioned slurry160. The resin-in-pulp step 164 can be performed by any suitabletechnique with any suitable ion exchange resin. Examples of suitabletechniques include that discussed in U.S. patent application, Ser. No.09/452,736, filed in Jun., 2000, entitled “A Process for Recovering Goldfrom Thiosulfate Leach Solutions and Slurries with Ion Exchange Resins”,to Thomas, et al.; U.S. patent application Ser. No. 09/034,846, filedMar. 4, 1998, entitled “Method for Recovering Gold from RefractoryCarbonaceous Ores”; and U.S. Pat. Nos. 5,536,297 and 5,785,736, all ofwhich are incorporated herein by reference. Preferred resins includeanion exchange resins, preferably a strong base resin including aquaternary amine attached to a polymer backbone. A strong base resin ispreferred over a weak base resin. The precious metal loading capacity ofa strong base resin is typically greater than that of a weak base resin,such that a lower volume of resin is required. Gel resins andmacroporous resins are suitable. Suitable resins include all commercialstrong-base resins of either Type I (triethylamine functional groups) orType II (triethyl ethanolamine functional groups). Specific strong-baseion exchange resins include “A500” manufactured by Purolite, “A600”manufactured by Purolite, “21K” manufactured by Dow Chemical, “AmberliteIRA 410” manufactured by Rohm and Haas, “Amberlite IRA 900” manufacturedby Rohm and Haas, and “Vitrokele 911 ” supplied by Signet. Because theRIP pretreatment and resin-in-pulp steps 140 and 164 are preferablyperformed in the same vessel (though they may be performed in differentvessels), the temperature, leach slurry pH, and residence time typicallydepend on which of the above techniques are used to reduce thepolythionate concentration.

[0079] Resin-in-pulp treatment can be performed in any suitable vessel.A preferred vessel is a Pachuca tank, which is an air-agitated, conicalbottomed vessel, with air being injected at the bottom of the cone. Anadvantage of the Pachuca system is reduced resin bead breakage andimproved dispersion of the resin beads in the slurry as compared tomechanically agitated systems. The RIP recovery is preferably carriedout in four or more tanks connected in series, more preferably betweenfour and eight such Pachuca tanks.

[0080] During resin-in-pulp 164, the resin will become “loaded” with thedissolved precious metals. Typically, at least about 99% and moretypically at least about 99.8% of the precious metal(s) in the leachslurry will be “loaded” or adsorbed onto the resin.

[0081] To inhibit the formation ofpolythionates and the consequentprecious metal recovery problems and increased reagent consumption, theleach slurry can be maintained in an inert (or an at least substantiallynonoxidizing) atmosphere and/or an inert (or an at least substantiallynonoxidizing) gas can be sparged through the leach slurry. Theatmosphere is preferably maintained (and/or gas sparging used) duringRIP pretreatment 140 and resin-in-pulp 164. As used herein, “inert”refers to any gas which is at least substantially free of oxidants, suchas molecular oxygen, that can cause thiosulfate to be converted into apolythionate. For example, an “inert” gas would include a reducing gas.Typically, the inert atmosphere will include at least about 85 vol % ofan inert gas, preferably nitrogen gas, and no more than about 5 vol %oxidants, such as oxygen gas, that can cause thiosulfate conversion intoa polythionate. The molecular nitrogen can be a byproduct of the oxygenplant that is employed in the leaching step to provide superatmosphericpartial pressures of oxygen gas. As will be appreciated, the leachslurry 144 during transportation between the leaching and RIPpretreatment steps 132 and 140 and if applicable from the RIPpretreatment and resin-in-pulp steps 140 and 164 (except duringconditioning 156) is typically in a conduit that is not open to thesurrounding atmosphere. If the leach slurry is open to an atmosphereduring transportation in either or both of these stages, the leachslurry should be maintained in the presence of the inert atmosphereduring any such transportation.

[0082] While not wishing to be bound, it is believed that sparging ismore effective than an inert atmosphere without sparging in controllingpolythionate production. Sparging appears to inhibit molecular oxygeningress into the solution, even where the reactor is open to the ambientatmosphere, because of the outflow of inert gas from the surface of thesolution.

[0083] The barren leach slurry 168 (which will typically contain no morethan about 0.01 ppm precious metals or 1% of the precious metal(s) inthe leach solution 144) is subjected to one or more stages of countercurrent decantation (“CCD”) 172. In CCD 172, the solids are separated inthe underflow 176 from the barren leach (or overflow) solution 180 andsent to the tailings pond. The barren leach solution 180 is separated inthe overflow from the solids and forwarded to regeneration step 184 toconvert polythionates to thiosulfate. As will be appreciated, CCDperforms liquid/solid separation, provides water balancing in thecircuit, and prevents build up of impurities in the leach circuit byremoving a portion of the leach solution with the solids.

[0084] Regeneration 184 can be performed in one or more vessel(s) and/orby in line sulfide (and/or sulfite) addition to a conduit carrying thestripped lixiviant solution. If a number of the techniques are employed,they can be performed simultaneously (in the same reactors) orsequentially (in different reactors), as desired.

[0085] The regenerated lixiviant solution 128 is recycled to thegrinding step 108 along with the thickener overflow 120 and ultimatelyto the leaching step 132.

[0086] The loaded resin 188 is screened 190 and washed with water toremove any leach slurry (liquid and/or leached material) from the resinbeads.

[0087] The recovered beads 192 are contacted with an eluant to strip orelute 194 adsorbed precious metal into the eluate and form a pregnantsolution 196 containing typically at least most (and more typically atleast about 95%) of the precious metal on the resin and a stripped resin197.

[0088] The eluant can be any suitable eluant that can displace theadsorbed precious metal from the loaded resin beads. The eluant couldinclude salts, such as one or more types of polythionate ions as setforth in U.S. application Ser. No. 09/452,736 above, and a nitrate, athiocyanate, a sulfite, a thiourea, a perchlorate and mixtures thereof.

[0089] Typically, the concentration of the eluant in the pregnantsolution 196 ranges from about 0.25 to about 3 M; the temperature ofelution 194 from about 5 to about 70° C., and the pH of elution 194 fromabout pH 5 to about pH 12. Under the conditions, at least about 90% andmore typically from about 95 to about 99% of the precious metal adsorbedon the resin is displaced by the eluant into the pregnant solution 196.

[0090] The stripped resin 197 is recycled to the resin-in-pulp step 164.Optionally, the stripped resin 197 can be regenerated (not shown) byknown techniques prior to reuse of the resin. As will be appreciated,the resin can be regenerated by acid washing the resin with an acid suchas nitric acid or hydrochloric acid. The acid wash removes adsorbedeluant and/or impurities from the resin and frees up the functionalsites on the resin surface (previously occupied by the eluant) to adsorbadditional precious metal. In the case of a polythionate eluant, theresin can be regenerated by contacting the resin with sulfide and/orsulfite to reduce the polythionate ions to thiosulfate ions and sulfateions. After regeneration, the resin and regeneration product solutionare separated by screening and washing.

[0091] The pregnant solution 196, which includes the eluant andtypically no more than about 100 ppm and more typically from about 10 toabout 500 ppm solubilized precious metals, is subjected toelectrowinning 198 to recover the solubilized precious metals and form abarren solution 199. Problems in electrowinning of precious metals outof a medium containing polythionates and/or thiosulfate have beenencountered in U.S. patent application Ser. No. 09/452,736. When theeluant is a polythionate the polythionate and thiosulfate tend to beco-reduced with the precious metal at the cathode to produce elementalsulfur, which interferes with the efficient continued operation of theelectrowinning circuit while the polythionate and thiosulfate are alsowastefully oxidized to sulfate ions at the anode.

[0092] These problems are overcome by the present invention through theuse of sulfite in the pregnant solution. Sulfite is added to the eluantand/or to the pregnant solution 196 prior to, during, or afterelectrowinning. Preferably, sulfite is added to the eluant prior to theelution step 194. In the presence of sulfite, the precious metal isreduced at the cathode while the sulfite is oxidized to sulfate at theanode. This has the benefit of lowering the cell voltage required.Preferably, the concentration of sulfite in the pregnant solution 196(in the elution and electrowinning steps 194, 198) is at least about0.01M and more preferably ranges from about 0.1 to about 2 M. Thesulfite is preferably in the pregnant solution with another eluant, suchas any of the eluants noted above.

[0093] The stripped or barren solution 199 is removed from theelectrowinning cell(s) and returned to the elution step 194. Ableedstream (not shown) of the barren solution 199 can be used tocontrol buildup of impurities such as sulfate.

[0094] The recovered precious metal 195, which contains the preciousmetal recovered in electrowinning and impurities, is subjected toretorting 193 by known techniques to remove the impurities and formprecious metal sludge. The sludge, which contains at least most of theprecious metal in the recovered precious metal 195, is refined toproduce a precious metal product of high purity.

[0095]FIG. 2 depicts another embodiment of a process for thiosulfateleaching of a refractory precious metal-containing material. FIG. 2shows an alternative to resin-in-pulp for precious metal recovery.Following leaching 132, the precious metal bearing solution 144 isseparated 200 from the solids by any suitable means, such as bycounter-current decantation washing and/or filtration. Preferably, atleast about 95% and more preferably at least about 99% of the preciousmetal is separated from the solids with the latter going to tailingsimpoundment.

[0096] The separated precious metal bearing solution 204 is directed tothe precious metal precipitation—thiosulfate regeneration step 208. Thisprocess can be carried out in any suitably agitated reactor or pluralityof agitated reactors. The pH of the precious metal bearing solution 204is adjusted if necessary to about pH 5.5-12, more preferably about pH7-11, even more preferably about pH 9-11 using a suitable basic reagentsuch as sodium hydroxide and the solution is contacted with a reductant,preferably a sulfide and/or bisulfide and/or polysulfide reagent toprecipitate at least about 99% of the precious metal and convert atleast about 90% of the polythionates to thiosulfate, effectivelyregenerating the thiosulfate lixiviant. The effectiveness of theconversion causes significantly less thiosulfate reagent to be consumedduring the process than for conventional thiosulfate leaching processes.The use of a sulfide and/or bisulfide and/or polysulfide has the addedbenefit of reducing impurities such as copper or mercury or manganesefrom solution thereby reducing the rate of thiosulfate degradation.While not wishing to be bound by any theory, it is believed that themost likely composition of the precipitate is the metallic preciousmetal and/or a precious metal sulfide, such as Au₂S Maximumprecipitation of gold and regeneration of thiosulfate is accomplished byadding at least a stoichiometric amount of reductant (relative to thedissolved precious metal and polythionate concentrations) to reduce thesolution ORP to at least about −150 mV (SHE). The temperature ispreferably maintained in the range of about 5 to 40° C., and morepreferably at ambient temperature, about 20° C. The retention time isabout 5 minutes to about 2 hours, more preferably about 15 minutes toabout 1 hour. The process is conducted under oxygen-depleted conditions,with the solution preferably containing no more than about 1 ppmdissolved molecular oxygen and more preferably less than about 0.2 ppmdissolved molecular oxygen concentration, by bubbling anoxygen-deficient gas such as nitrogen into the slurry and/or maintaininga blanket of nitrogen in the atmosphere over the slurry as noted above.

[0097] The precious metal bearing precipitate is separated from theregenerated solution 212 by any suitable method such as filtration, CCD,and the like and the separated precious metal 216 is recovered byrefining in furnaces.

[0098] The regenerated solution 220 is directed to the conditioning step224, which can be conducted in any suitably agitated reactor orplurality of reactors. The solution pH is adjusted to a value suitablefor recycling the solution back to grinding 108 and/or for preciousmetal scavenging 228. Preferably, the pH ranges from about pH 7 to aboutpH 12, more preferably about pH 8 to pH 10. The solution 220 is agitatedin the presence of an oxygen-containing atmosphere, such as air, tooxidize any remaining reductant (such as sulfide or bisulfide orpolysulfide) carried over from the precious metalprecipitation—thiosulfate regeneration step 208. The duration of theconditioning step 224 is preferably not sufficient to cause more thanabout 5% of the thiosulfate to form polythionates, or to yield apolythionate concentration of more than about 0.003M. The majority(typically at least about 80 vol %) of the conditioned solution 232 isthen recycled in recycle solution 236.

[0099] A minor portion (e.g., from about 2 to about 20 vol %) of theconditioned solution or bleed stream 240 may have to be bled to tailingsto control the buildup of impurities, such as soluble sulfate andmetallic impurities. Prior to discharge to tailings the bleed portion240 of the conditioned solution 232 is directed to the precious metalscavenging step 228 to recover any precious metals remaining in solutionthat were not recovered in the precious metal precipitation—thiosulfateregeneration step 208. Precious metal scavenging can be accomplished, byany suitable gold recovery technique such as by passing the bleedsolution 240 through a column containing a strong base resin to adsorbthe precious metal. While not wishing to be bound by any theory,precipitated precious metal can be redissolved due to trace amount ofmolecular oxygen in the solution and incomplete reductionofpolythionates in the solution. Because the amount of polythionates inthe bleed is negligible, a resin-in-column recovery technique will havean excellent ability to load any remaining dissolved precious metal.

[0100] In an alternative configuration (not shown), the precious metalprecipitates are redissolved in a suitable solvent, such asnitric/hydrochloric acid, cyanide, thiosulfate, thioureachloride/chlorine and bromide/bromine to provide a preciousmetal-containing solution. The precious metal can then be recovered byelectrolysis as noted above in connection with step 198 of FIG. 1.

[0101] This process is preferred in certain applications over theprocess of FIG. 1. For certain precious metal-containing materials, itis difficult to obtain high rates of precious metal adsorption ontoresins while maintaining the precious metal in solution. The use of anRIP pretreatment step, though beneficial, can be difficult to usewithout experiencing some precious metal precipitation. Conditioning 156may not be completely effective in redissolving gold precipitates, whichwould be discarded with the barren solids to tailings. The process ofFIG. 2 can also be more robust, simpler, and therefore easier to designand operate than the process of FIG. 1.

[0102]FIG. 3 shows an alternative to FIG. 2 in which thiosulfateleaching is conducted in two stages to achieve more effective recoveryof the precious metal content. Leaching is first conducted atatmospheric pressure and ambient temperature in the presence of anoxygen-containing gas such as air or industrially available oxygen (step300) to dissolve from about 30 to 95% of the leachable precious metalcontent. The leachable precious metal content is defined as that portionof the precious metal content that is physically accessible to thethiosulfate lixiviant and is not encapsulated within constituentscontained in the host material. The precious metal bearing solution 304is separated from the solids 308 (step 200), the solids 308 are repulpedwith a portion 310 of the recycle solution 236, and the resulting slurry308 is then directed to pressure leaching (step 312) to dissolve themajority, ie. about 5-70%, of the remaining leachable precious metalcontent that was not recovered in atmospheric leaching 300. In pressureleaching the solids are leached under superatmospheric conditions suchas the conditions described previously (step 132 of FIG. 1). Themolecular oxygen partial pressure in leach 300 preferably ranges fromthe molecular oxygen partial pressure at ambient conditions (e.g., morethan about 3 psia at sea level) to about 15 psia and the molecularoxygen partial pressure in leach 312 preferably ranges from more than 15psia to about 500 psia. The slurry 316 exiting pressure leaching 312 isthen processed in essentially the same manner as the slurry exitingleaching 300 in FIG. 2. That is, the slurry 316 is subjected tosolid/liquid separation 320 in the presence of wash water to separatethe barren solid material 324 from the (second) pregnant leach solution328. The first and second pregnant leach solutions 304, 328 aresubjected to precious metal precipitation—thiosulfate regeneration 208,further solid/liquid separation 212, conditioning 224 and precious metalscavenging 228 as noted above in connection with FIG. 2.

[0103] The process of FIG. 3 typically performs the bulk of theleaching, or precious metal dissolution, under ambient conditions, whichis much cheaper than leaching under superatmospheric conditions. Themore-difficult-to-dissolve precious metals and weakly preg-robbedprecious metals are then dissolved in a higher pressure leach. Becauseless precious metal remains to be dissolved, the high pressure leach canhave a shorter residence time and therefore lower capacity than would bepossible in the absence of the ambient pressure leach.

[0104]FIG. 4 depicts another embodiment of the present invention. Theprocess is similar to those discussed above except that thiosulfateleaching is performed by heap leaching 400 techniques. The comminutedprecious metal-containing material 404 can be directly formed into aheap (in which case the material would have a preferred P80 size of fromabout 2 inches to about {fraction (1/4)} inch, possibly agglomerated andformed into a heap. The thiosulfate lixiviant (which commonly includes arecycled thiosulfate lixiviant 236 mixed with a makeup (fresh)thiosulfate solution(not shown)) is applied to the top of the heap usingconventional techniques, and the pregnant leach solution 408 iscollected from the base of the heap. Refining can be performed using anyof the techniques noted above.

[0105] To facilitate extraction of gold from sulfidic and/orcarbonaceous materials, the thiosulfate leach step in any of the aboveprocesses can be preceded by one or more pretreatment steps to destroyor neutralize the carbon-containing and/or sulfidic minerals. By way ofexample, the intermediate steps can include one or more of biooxidationor chemical oxidation to oxidize sulfides, ultrafine grinding toliberate occluded precious metals, conventional roasting to destroycarbon- and/or sulfide-containing minerals, and/or microwave roasting.

EXAMPLE 1

[0106] A gold ore from Nevada, designated Sample A, was subjected tothiosulfate leaching under oxygen pressure at varying temperatures. Theore assayed 24.1 g/t gold, 2.59% iron, 0.31% total sulfur, 0.28% sulfidesulfur, 3.40% total carbon, 1.33% organic carbon and 0.02% graphiticcarbon. From a diagnostic leaching analysis of the ore it was determinedthat a maximum of 83% of the contained gold was capable of beingsolubilized while the remaining gold was inaccessible to a lixiviantbecause it was encapsulated within pyrite and/or other mineralscontained in the ore.

[0107] The ore was ground to 80% passing 200 mesh (75 μm). Samples ofthe ore were slurried with water to a pulp density of 33% solids in amechanically agitated laboratory autoclave. The natural pH of the slurryat ambient temperature was 8.3. The pH of the slurry was adjusted to 9with sodium hydroxide and a quantity of sodium thiosulfate reagent wasadded to adjust the initial leach solution thiosulfate concentration to0.1 molar (M). The autoclave was sealed and pressurized to 100 psigoxygen with pure (95% plus) oxygen gas and the slurry was heated to thedesired temperature (if required). Leaching was maintained for 6 hours,during which pulp samples were taken at 2 and 4 hours in order tomonitor gold extraction with time. Upon termination of leaching, theslurry was filtered and the residue solids were washed with a dilutethiosulfate solution. The residue solids and leach solution were assayedfor gold to determine the final gold extraction.

[0108] The results were as follows: Leach Temp. Leach Time Calc'd HeadResidue Au Ext'n (° C.) (hours) Au (g/t) Au (g/t) (%) 20 2 33.3 4 41.9 622.8 9.44 58.5 40 2 51.2 4 55.1 6 26.4 9.25 64.9 60 2 63.7 4 68.5 6 22.84.26 81.3 60 (repeat) 2 65.2 4 73.0 6 80.9

[0109] The results indicate that the rate and extent of gold extractionwas improved with increasing temperature and leach time in thetemperature range 20-60° C. The best results were obtained at 60° C.,with about 8 1% gold extraction obtained after 6 hours leaching, thisrepresenting about 98% of the leachable gold content of the ore.

EXAMPLE 2

[0110] A second gold ore from Nevada, designated Sample B, was subjectedto thiosulfate leaching under oxygen pressure at varying initial pH's.The ore assayed 9.45 g/t gold, 2.50% iron, 0.39% total sulfur, 0.36%sulfide sulfur, 4.20% total carbon, 1.46% organic carbon and 0.05%graphitic carbon. From a diagnostic leaching analysis of the ore it wasdetermined that 82% of the contained gold was capable of beingsolubilized. Samples of the ore were prepared and leached as describedin Example 1, except the temperature was 60° C. in each test, theautoclave was pressurized with 50 psig oxygen, the initial pH wasadjusted to either 9, 11 or 12, and the leach retention time wasextended to 8 hours for the pH 11 and 12 tests.

[0111] The results were as follows: Initial Leach Time Calc'd HeadResidue Au Ext'n pH (hours) Au (g/t) Au (g/t) (%) 9 1 50.2 2 62.4 4 72.06 8.49 2.10 75.3 11 1 41.3 2 63.0 4 69.3 8 8.61 2.00 76.8 12 1 6.4 2 1.04 13.6 8 8.61 3.34 61.2

[0112] The results indicate that there was not much difference in goldleaching behaviour over the initial pH range of 9-11 (it should be notedthat the pH tended to decline during leaching). However, gold leachingwas suppressed during the first 4 hours of leaching at pH 12, but thenstarted to recover.

EXAMPLE 3

[0113] A third gold ore sample from Nevada, Sample C, was subjected tothiosulfate leaching under oxygen pressure at varying temperatures. Thehead analysis of the ore was as follows: Gold Ore Sample C Au, g/t 9.50C (t), % 4.45 Fe, % 2.52 C (CO₃), % 3.12 Cu, ppm 40 C (org), % 1.38 As,ppm 647 S (2−), % 0.35 Hg, ppm 14 S (t), % 0.27 Ca, % 9.0 Mg, % 1.5

[0114] From a diagnostic leaching analysis of the ore it was determinedthat 83% of the contained gold was capable of being solubilized.

[0115] The ore was ground to 80% passing 200 mesh (75 μm). Samples ofthe ore were slurried with water to a pulp density of 33% solids in amechanically agitated laboratory autoclave. The initial pH of the slurrywas adjusted to approximately 11 with sodium hydroxide, after which theautoclave was sealed and pressurized to 100 psig oxygen with pure (95%plus) oxygen gas and the slurry was heated to the desired temperature.To initiate leaching, a quantity of sodium thiosulfate stock solutionwas injected to adjust the leach solution thiosulfate concentration to0.1 M. Leaching was continued for 6 to 10 hours, during which noadditional reagents were added. Pulp samples were taken at set intervalsduring leaching in order to monitor gold extraction with time. Upontermination of leaching, the slurry was filtered and the residue solidswere washed with a dilute thiosulfate solution. The residue solids andleach solution were assayed for gold to determine the final goldextraction.

[0116]FIG. 5 depicts graphically the effect of leach temperature, in therange 40-80° C., on the rate of gold extraction from Sample C. It can beseen that the gold leached quickly at 60° C. and 80° C., there beinglittle difference in the extraction rate at the two temperatures. Thegold extraction peaked at approximately 83%, the maximum extractable,after 6 hours leaching. Gold leaching was slowed if the temperature waslowered to 40° C., but 80% gold extraction was still obtained after 10hours leaching at 40° C.

[0117] An overall summary of the results is provided below: Test #6 Test#25 Test #15 Parameter 80° C. 60° C. 40° C. Leach time, hours 8 6 10Final pH 7.0 8.7 9.3 Final ORP, mV (SHE) 307 242 225 Calc'd Head Au, g/t9.48 9.43 9.27 Residue Au, g/t 1.59 1.63 1.81 Au Ext'n, % 83.2 82.7 80.5

EXAMPLE 4

[0118] The gold ore designated Sample C was subjected to thiosulfateleaching at varying oxygen pressures. Samples of the ore were preparedand leached as described in Example 3 except the temperature wasmaintained at 60° C. in each test and the oxygen partial pressure wasvaried.

[0119]FIG. 6 portrays the effect of oxygen partial pressure, in therange 0-200 psig, on the rate of gold extraction from Sample C (in the 0psig O₂ test, the autoclave was not pressurized but the head space wasmaintained with pure oxygen at atmospheric pressure). It can be seenthat the rate of gold leaching was somewhat sensitive to oxygenpressure, in that the rate increased with increasing pressure,particularly during the first two hours of leaching. After 6 hoursleaching, gold extraction varied from a low of 78% at 0 psig O₂ to ahigh of 83% at 200 psig O₂.

[0120] An overall summary of the results is provided below: Test #7 Test#25 Test #22 Test #28 Test #31 Parameter 200 psig O₂ 100 psig O₂ 50 psigO₂ 10 psig O₂ 0 psig O₂ Leach time, hours 8 6 6 6 6 Final pH NA 8.7 9.09.3 9.4 Final ORP, mV (SHE) NA 242 223 216 232 Calc'd Head Au, g/t 9.789.43 9.40 8.95 9.08 Residue Au, g/t 1.68 1.63 1.77 1.72 2.00 Au Ext'n, %82.8 82.7 81.1 80.8 78.0

EXAMPLE 5

[0121] The gold ore designated Sample C was subjected to thiosulfateleaching under oxygen pressure at varying initial sodium thiosulfateconcentrations. Samples of the ore were prepared and leached asdescribed in Example 3 except the temperature was maintained at 60° C.in each test and the initial sodium thiosulfate concentration wasvaried.

[0122]FIG. 7 portrays the effect of initial sodium thiosulfateconcentration, in the range 0.05-0.2 M, on the rate of gold extractionfrom Sample C. It can be seen that the rate of gold leaching wasinsensitive to initial thiosulfate concentration in the 0.1-0.2 M range.At 0.05 M thiosulfate, the rate was reduced significantly, particularlyduring the first two hours of leaching. After 6 hours leaching goldextraction was 78% at 0.05 M thiosulfate compared to 82% achieved atboth 0.1 M and 0.2 M thiosulfate concentration.

[0123] An overall summary of the results is provided below: Test #4 Test#25 Test #8 Parameter 0.2 M 0.1 M 0.05 M Leach time, hours 8 6 6 FinalpH 8.7 8.7 8.5 Final ORP, mV (SHE) NA 242 262 Calc'd Head Au, g/t 8.859.43 9.40 Residue Au, g/t 1.50 1.63 1.87 Au Ext'n, % 83.0 82.7 80.1

EXAMPLE 6

[0124] The gold ore designated Sample C was subjected to thiosulfateleaching under oxygen pressure at two different pulp densities. Samplesof the ore were prepared and leached as described in Example 3, exceptthe temperature was maintained at 60° C. in each test and the leach pulpdensity was either 33% or 45% solids by weight.

[0125]FIG. 8 portrays the effect of 33% vs. 45% pulp density on the rateof gold extraction from Sample C. The rate of gold leaching was found tobe essentially insensitive to pulp density in this range.

[0126] An overall summary of the results is provided below: Test #26Test #25 45% pulp 33% pulp Parameter density density Leach time, hours 66 Final pH 8.5 8.7 Final ORP, mV (SHE) 286 242 Calc'd Head Au, g/t 9.299.43 Residue Au, g/t 1.68 1.63 Au Ext'n, % 81.9 82.7

EXAMPLE 7

[0127] A fourth gold ore sample from Nevada, Sample D, was subjected tothiosulfate leaching at 60° C. and 10 psig oxygen pressure at thenatural pH of the slurry, for 8 hours. The head analysis of the ore wasas follows: Gold Ore Sample D Au, g/t 12.15 C (t), % 4.31 Fe, % 2.09 C(CO₃), % 3.02 Cu, ppm 39 C (org), % 1.30 As, ppm 692 S (2−), % 0.12 Hg,ppm 27 S (t), % 0.22 Ca, % 8.9 Mg, % 1.3

[0128] From a diagnostic leaching analysis of the ore it was determinedthat 80% of the contained gold was capable of being solubilized.

[0129] The ore was ground to 80% passing 200 mesh (75 μm). A sample ofthe ore was slurried with water to a pulp density of 40% solids in amechanically agitated laboratory autoclave. The autoclave was sealed andpressurized to 100 psig oxygen with pure (95% plus) oxygen gas and theslurry was heated to 60° C. To initiate leaching, a quantity of sodiumthiosulfate stock solution was injected to adjust the leach solutionthiosulfate concentration to 0.1 M. Leaching was continued for 8 hours,during which no additional reagents were added. Pulp samples were takenat set intervals during leaching in order to monitor gold extraction andremaining thiosulfate with time. Upon termination of leaching, theslurry was filtered and the residue solids were washed with a dilutethiosulfate solution. The residue solids and leach solution were assayedfor gold to determine the final gold extraction.

[0130]FIG. 9 depicts percent gold extraction and percent remainingthiosulfate with time. Gold extraction reached 79.3% after 8 hours, or99% of the leachable gold content. Thiosulfate consumption was low, with86.7% of the thiosulfate remaining after 8 hours and available forreuse.

[0131] An overall summary of the results is provided below: ParameterTest #37-01 Leach time, hours 8 Initial pH 7.9 Final pH 9.0 Initial ORP,mV (SHE) 411 Final ORP, mV (SHE) 397 Calc'd head Au, g/t 11.50 ResidueAu, g/t 2.38 Gold extraction, % 79.3 Amount of thiosulfate remaining, %86.7

EXAMPLE 8

[0132] A thiosulfate leach discharge slurry was heated to 60° C. in anagitated reactor in preparation for RIP pre-treatment, the objectivebeing to reduce the polythionate content without precipitating gold. Theslurry was kept under a nitrogen atmosphere to ensure the dissolvedoxygen content was maintained below 0.2 mg/L. A single dose of a 0.26 Msodium bisulfide (NaHS) solution, adjusted to pH 9, was added and thepretreatment was allowed to proceed at 60° C. and ambient pressure for 2hours. The amount of sulfide added was 150% of stoichiometric based onthe amount required to convert the polythionates back to thiosulfate inaccordance with the following reactions:

2S₄O₆ ²⁻+S²⁻+{fraction (3/2)}H₂O→{fraction (9/2)}S₂O₃ ²⁻+3H⁺

S₃O₆ ²⁻+S²⁻→2S₂O₃ ²⁻

[0133] A summary of the results is provided below: Time Au S₂O₃ ²⁻ S₄O₆²⁻ S₃O₆ ²⁻ ORP (min) (mg/L) (g/L) (g/L) (g/L) (mV) pH 0 4.36 8.38 0.510.59 240 6.9 120 4.36 11.0 0.06 0.10 5 6.7

[0134] The tetrathionate and trithionate concentrations were reduced by88% and 83% respectively while all of the gold remained in solution.

EXAMPLE 9

[0135] A pregnant thiosulfate leach solution was adjusted to pH 10 withsodium hydroxide in an agitated reactor in preparation for sulfidetreatment, the objective being to regenerate thiosulfate and precipitatethe gold. The solution was kept under a nitrogen atmosphere to ensurethe dissolved oxygen content was maintained below 0.2 mg/L. A singledose of a 0.26 M sodium sulfide (Na₂S) solution was added and thetreatment was allowed to proceed for 2 hours at ambient temperature (22°C.) and pressure. The amount of sulfide added was 100% of stoichiometricbased on the amount required to convert the polythionates back tothiosulfate in accordance with the following reactions:

2S₄O₆ ²⁻+S²⁻+{fraction (3/2)}H₂O→{fraction (9/2 )}S₂O₃ ²⁻+3H⁺

S₃O₆ ²⁻+S²⁻→2S₂O₃ ²⁻

[0136] A summary of the results is provided below: Time Au S₂O₃ ²⁻S₄O₆2⁻ S₃O₆ ²⁻ ORP (min) (mg/L) (g/L) (g/L) (g/L) (mV) pH 0 4.12 7.80.84 1.47 200 10.0 10 0.05 9.9 0.01 0.01 −210 11.0 20 0.02 9.9 0.01 0.01−220 10.4 30 0.01 9.9 0.01 0.01 −230 10.2 60 0.01 9.8 0.01 0.01 −26010.3 90 0.01 10.1 0.01 0.01 −260 10.3 120 0.01 10.2 0.01 0.01 −260 10.3

[0137] The rate of conversion of polythionates to thiosulfate wasextremely fast under ambient conditions, with essentially completeconversion achieved after about 10 minutes. Similarly, goldprecipitation was also fast and essentially complete after about 30minutes.

[0138] While this invention has been described in conjunction with thespecific embodiments thereof, it is evident that many alternatives,modifications, and variations will be apparent to those skilled in theart. Accordingly, preferred embodiments of the invention as set forthherein are intended to be illustrative, not limiting. By way of example,any source of sulfur species with an oxidation state less than +2 may beused in any of the above process steps to convert polythionates tothiosulfate. The regeneration step 184 in FIG. 1 can be performed in avariety of locations. For example, regeneration 184 can be performed inthe recycle loop after CCD 172 and before grinding 108, between grinding108 and thickening 116, in the thickener 116 immediately before orduring, leaching 132 and/or between resin in pulp 164 and CCD 172. Freshthiosulfate can also be added in a number of locations. For example,fresh thiosulfate can be added in any of the locations referred topreviously for the regeneration step 184 and can be added after orduring regeneration 184 as noted above or in a separate tank orlocation. In FIG. 3, a lixiviant other than thiosulfate, such ascyanide, can be used in the atmospheric leach 300 with thiosulfate beingused in the pressure leach 312. These and other changes may be madewithout departing from the spirit and scope of the present invention.

What is claimed is:
 1. A process for recovering a precious metal from aprecious metal-containing material, comprising: contacting the preciousmetal-containing material with a thiosulfate lixiviant atsuperatmospheric pressure in the absence of at least one of added copperand added ammonia to solubilize the precious metal and form a pregnantthiosulfate leach solution containing the solubilized precious metal;and thereafter recovering the solubilized precious metal from thepregnant thiosulfate leach solution.
 2. The process of claim 1, whereinin the contacting step the thio sulfate lixiviant has a pH of less than9 and a temperature of from about 40 to about 80° C., and the totalammonia concentration in the thiosulfate lixiviant is less than 0.05M.3. The process of claim 1, wherein in the contacting step thethiosulfate lixiviant contains less than 20 ppm of copper ion.
 4. Theprocess of claim 1, wherein the lixiviant includes dissolved molecularoxygen.
 5. The process of claim 1, wherein the thio sulfate lixiviantcontains no more than about 0.01 M of added sulfite.
 6. The process ofclaim 1, wherein the thiosulfate lixiviant includes sulfite and a totalconcentration of sulfate in the thiosulfate lixiviant is no more thanabout 0.02M.
 7. A precious metal recovered by the process of claim
 1. 8.A process for recovering a precious metal from a carbonaceous preciousmetal-containing material, comprising: contacting the carbonaceousprecious metal-containing material with a thiosulfate lixiviant in thesubstantial or complete absence of added copper and added ammonia tosolubilize the precious metal and form a pregnant thiosulfate leachsolution containing the solubilized precious metal; and thereafterrecovering the solubilized precious metal from the pregnant thiosulfateleach solution.
 9. The process of claim 8, wherein in the contactingstep the thiosulfate lixiviant has a pH of less than 9 and a temperatureof from about 40 to about 80° C., and the total ammonia concentration inthe thiosulfate lixiviant is less than about 0.05M.
 10. The process ofclaim 8, wherein in the contacting step the thiosulfate lixiviantcontains less than 20 ppm of copper ion.
 11. The process of claim 8,wherein the lixiviant includes molecular oxygen.
 12. The process ofclaim 8, wherein the thiosulfate lixiviant contains no more than about0.01M of added sulfite.
 13. The process of claim 8, wherein thethiosulfate lixiviant includes sulfite and a total concentration ofsulfite in the thiosulfate lixiviant is no more than about 0.02M.
 14. Aprecious metal recovered by the process of claim
 8. 15. A process forrecovering a precious metal from a precious metal-containing material,comprising: (a) contacting the precious metal-containing material with athiosulfate lixiviant to solubilize the precious metal and form apregnant thiosulfate leach solution containing the solubilized preciousmetal and a polythionate; (b) contacting the pregnant leach solutionwith a sulfide-containing reagent to precipitate at least most of thesolubilized precious metal and convert at least most of the polythionateto thiosulfate; and (c) thereafter recovering the precious metalprecipitate from the thiosulfate leach solution.
 16. The process ofclaim 15, further comprising before the contacting step (b): separatingat least most of the precious metal-containing material from at leastmost of the pregnant thiosulfate leach solution.
 17. The process ofclaim 15, wherein the sulfide-containing reagent is at least one of apolysulfide other than a bisulfide, a bisulfide, and a sulfide otherthan a bisulfide and a polysulfide.
 18. The process of claim 15, whereinthe pregnant leach solution in contacting step (b) has a pH ranging fromabout pH 5.5 to about pH
 12. 19. The process of claim 15, wherein in thecontacting step (b), the pregnant leach solution has a dissolvedmolecular oxygen content of no more than about 1 ppm.
 20. The process ofclaim 15, wherein the thereafter recovering step includes separating theprecious metal precipitates from a barren leach solution.
 21. Theprocess of claim 20, further comprising: (d) adjusting a pH of thebarren leach solution to a pH of from about pH 7 to about pH 12; (e)contacting the barren leach solution with a gas including at least about5 vol. % molecular oxygen to oxidize any remaining sulfide-containingreagent; and (f) scavenging precious metal from at least a portion ofthe barren leach solution.
 22. A process for recovering a precious metalfrom a precious metal-containing material, comprising: (a) solubilizinga first portion of the precious metal in the precious metal-containingmaterial to form a first pregnant leach solution, wherein thesolubilizing step (a) is conducted at a first oxygen partial pressure;(b) solubilizing a second portion of the precious metal in the preciousmetal-containing material to form a second pregnant leach solution,wherein the solubilizing step (b) is conducted at a second oxygenpartial pressure and wherein the first oxygen partial pressure is lessthan the second oxygen partial pressure; (c) separating at least thesecond pregnant leach solution from the precious metal-containingmaterial; and (d) recovering the solubilized precious metal from thefirst and second pregnant leach solution.
 23. The process of claim 22,wherein the first pregnant leach solution is separated from the preciousmetal-containing material before solubilizing step (b).
 24. The processof claim 22, wherein a first lixiviant in step (a) is different from asecond lixiviant in step (b).
 25. The process of claim 22, wherein inboth steps (a) and (b) the precious metal is solubilized in athiosulfate lixiviant.
 26. A process for recovering a precious metalfrom a precious metal-containing material, comprising: contacting theprecious metal-containing material with a thiosulfate lixiviant atsuperatmospheric pressure and at a pH less than pH 9 to solubilize theprecious metal and form a pregnant thiosulfate leach solution containingthe solubilized precious metal; and recovering the solubilized preciousmetal from the pregnant thiosulfate leach solution.
 27. The process ofclaim 26, wherein at least most of the precious metal in the preciousmetal-containing material is solubilized in the contacting step.
 28. Theprocess of claim 26, wherein the precious metal-containing materialincludes carbonaceous minerals.
 29. The process of claim 28, wherein theprecious metal-containing material is a double refractory ore.
 30. Theprocess of claim 26, wherein the pH is less than about pH8.
 31. Theprocess of claim 26, wherein the partial pressure of molecular oxygenranges from about 4 to about 500 psia.
 32. The process of claim 26,wherein the total pressure in the contacting step ranges from about 15to about 600 psia.
 33. The process of claim 26, wherein the thiosulfatelixiviant is at least substantially free of ammonia.
 34. The process ofclaim 26, wherein the thiosulfate lixiviant includes no more than about20 ppm copper ion.
 35. The process of claim 26, wherein the thiosulfatelixiviant includes no more than about 0.02M sulfite.
 36. The process ofclaim 26, wherein the temperature in the contacting step ranges fromabout 5 to about 150° C.
 37. A precious metal recovered by the processof claim
 26. 38. A process for recovering a precious metal from aprecious metal-containing material, comprising: contacting the preciousmetal-containing material with a thiosulfate lixiviant atsuperatmospheric pressure and at a temperature ranging from about 40 toabout 80° C. to solubilize at least most of the precious metal in thematerial and form a pregnant thiosulfate leach solution containing thesolubilized precious metal; and recovering the solubilized preciousmetal from the pregnant thiosulfate leach solution.
 39. The process ofclaim 38, wherein the temperature is more than about 60° C.
 40. Theprocess of claim 38, wherein the molecular oxygen partial pressure inthe contacting step ranges from about 4 to about 500 psia.
 41. Theprocess of claim 38, wherein the total pressure in the contacting stepranges from about 15 to about 600 psia.
 42. The process of claim 38,wherein the pH in the contacting step is less than pH9.
 43. The processof claim 38, wherein the thiosulfate lixiviant is at least substantiallyfree of ammonia.
 44. The process of claim 38, wherein the thiosulfatelixiviant includes no more than about 20 ppm added copper ion.
 45. Theprocess of claim 38, wherein the thiosulfate lixiviant includes no morethan about 0.01M added sulfite.
 46. A precious metal recovered by theprocess of claim
 38. 47. A process for recovering a precious metal froma precious metal-containing material, comprising: (a) contacting theprecious metal-containing material with a thiosulfate lixiviant tosolubilize the precious metal and form a pregnant thiosulfate leachsolution containing the solubilized precious metal; (b) contacting thepregnant thiosulfate leach solution with an extraction agent at atemperature of more than about 60° C. to recover the precious metal fromthe pregnant thiosulfate leach solution and convert trithionates in thepregnant thiosulfate leach solution into thiosulfate; and (c) recoveringthe precious metal from the extraction agent.
 48. The process of claim47, wherein the extraction agent is at least one of a resin and asolvent extraction reagent.
 49. The process of claim 48, wherein theextraction is an anion exchange resin.
 50. The process of claim 47,wherein the contacting step (a) is performed at superatmosphericpressure, a pH of less than pH 9, and a temperature of from about 40 toabout 80° C.
 51. The process of claim 47, wherein the contacting step(a) is performed in the substantial absence of ammonia.
 52. The processof claim 47, wherein the contacting step (a) is performed in thesubstantial absence of copper ion.
 53. A precious metal recovered by theprocess of claim
 47. 54. A process for recovering a precious metal froma precious metal-containing material, comprising: contacting theprecious metal-containing material with a thiosulfate lixiviant tosolubilize the precious metal and form a pregnant thiosulfate leachsolution comprising solubilized precious metal, thiosulfate, and atleast one of trithionate and tetrathionate; after the contacting step,converting at least most of the at least one of trithionate andtetrathionate in the pregnant thiosulfate leach solution intothiosulfate; and thereafter recovering the solubilized precious metalfrom the pregnant thiosulfate leach solution.
 55. The process of claim54, wherein the converting step includes: contacting the pregnantthiosulfate leach solution with at least one of a sulfite and a sulfide.56. The process of claim 54, wherein the converting step includes:heating the pregnant thiosulfate leach solution to a temperature of atleast about 60° C.
 57. The process of claim 54, wherein in thecontacting step the thiosulfate lixiviant has a pH of less than 9 and atemperature of from about 40 to about 80° C., the pressure issuperatmospheric, and the thiosulfate lixiviant is at leastsubstantially free of ammonia.
 58. The process of claim 54, wherein inthe contacting step the thiosulfate lixiviant is at least substantiallyfree of cupric ion.
 59. A precious metal recovered by the process ofclaim
 54. 60. A process for recovering a precious metal from a preciousmetal-containing material, comprising: contacting the preciousmetal-containing material with a lixiviant to solubilize the preciousmetal and form a pregnant leach solution containing the solubilizedprecious metal; and thereafter electrowinning the precious metal in thepresence of sulfite.
 61. The process of claim 60, wherein the lixiviantincludes thiosulfate and further comprising: contacting the pregnantleach solution with an extraction agent at a temperature of less thanabout 70° C. to collect the precious metal and convert polythionates inthe pregnant thiosulfate leach solution into thiosulfate; and thereafterremoving the precious metal from the extraction agent.
 62. The processof claim 61, wherein the contacting step includes: (a) contacting theprecious metal-containing material with a first lixiviant at a firstpressure to solubilize a first portion of the precious metal in theprecious metal-containing material; and (b) contacting the preciousmetal-containing material with a second lixiviant at a second pressuregreater than the first pressure to solubilize a second portion of theprecious metal in the precious metal-containing material.
 63. Theprocess of claim 62, wherein the first and second lixiviants eachinclude thiosulfate.
 64. The process of claim 60, wherein in thecontacting step the thiosulfate lixiviant has a temperature of fromabout 40 to about 80° C., the molecular oxygen partial pressure issuperatmospheric, and the lixiviant includes at least about 0.005Mthiosulfate and less than 0.05M of ammonia.
 65. The process of claim 60,wherein in the contacting step the thiosulfate lixiviant contains lessthan 20 ppm of added copper.
 66. A precious metal recovered by theprocess of claim
 60. 67. A process for recovering a precious metal froma precious metal-containing material, comprising: contacting theprecious metal-containing material with a thiosulfate leach solution tosolubilize the precious metal and form a pregnant thiosulfate leachsolution containing solubilized precious metal; maintaining a dissolvedmolecular oxygen content in at least one of the thiosulfate leachsolution and the pregnant thiosulfate leach solution at or below about 1ppm to inhibit the formation of trithionate and tetrathionate; andrecovering the solubilized precious metal from the pregnant thiosulfateleach solution.
 68. The process of claim 67, wherein, in the maintainingstep, the at least one of the thiosulfate leach solution and thepregnant thiosulfate leach solution is maintained in an atmosphere thatis at least substantially free of molecular oxygen, the atmosphere isinert and includes at least about 85 vol. % molecular nitrogen.
 69. Theprocess of claim 67, wherein the contacting step occurs in a reactor andthe pregnant leach solution is maintained in a molecular oxygen depletedatmosphere after removal from the reactor.
 70. The process of claim 67,wherein in the maintaining step a gas containing no more than about 5vol. % oxidants is sparged through the pregnant thiosulfate leachsolution.
 71. The process of claim 67, wherein in the contacting stepthe thiosulfate lixiviant contains less than 20 ppm of copper ion andless than 0.05M of ammonia.
 72. A precious metal recovered by theprocess of claim
 67. 73. A hydrometallurgical process for the recoveryof precious metal values from a refractory precious metal ore materialcontaining precious metal values and preg-robbing carbonaceouscompounds, comprising: (a) providing a body of particles and/orparticulates of the refractory precious metal ore material; (b)contacting the body of particles and/or particulates with a thiosulfatelixiviant solution at superatmospheric pressure and at a pH of less thanpH 9 to form stable precious metal thiosulfate complexes; (c) recoveringthe thiosulfate lixiviant solution from the body of particles and/orparticulates after a period of time which is sufficient for thethiosulfate lixiviant solution to become pregnant with precious metalvalues extracted from the ore material; and (d) recovering the preciousmetal values from the lixiviant solution.
 74. The process of claim 73,wherein in the contacting step (b) the thiosulfate lixiviant solutionhas a temperature of from about 40 to about 80° C., and contains lessthan 0.05M of ammonia, less than 20 ppm of added copper, and less thanabout 0.01M of added sulfite.
 75. The process of claim 73, wherein inthe recovering step (d) the dissolved precious metal is precipitatedfrom the lixiviant solution by the addition of at least one of a sulfideother than a polysulfide and a bisulfide, a bisulfide, and a polysulfideto the lixiviant solution.
 76. A precious metal recovered by the processof claim
 73. 77. A process for recovering a precious metal from aprecious metal-containing material, comprising: (a) contacting aprecious metal-containing material with a thiosulfate lixiviant todissolve the precious metal and form a pregnant thiosulfate leachsolution containing the dissolved precious metal; (b) contacting thepregnant thiosulfate leach solution with an adsorbent to load theprecious metal onto the adsorbent; (c) contacting the loaded adsorbentwith an eluant other than sulfite in the presence of sulfite to desorbthe precious metal adsorbed on the loaded adsorbent and form a loadedeluate containing the dissolved precious metal, and (d) recovering theprecious metal from the loaded eluate.
 78. The process of claim 77,wherein in the contacting step (c) the sulfite concentration ranges fromabout 0.01 to about 2M.
 79. The process of claim 77, wherein therecovering step includes electrowinning the precious from the loadedeluate.
 80. A process for recovering a precious metal for preciousmetal-containing material comprising: (a) contacting a preciousmetal-containing material with a thiosulfate lixiviant to dissolve theprecious metal and form a pregnant thiosulfate leach solution containingthe dissolved precious metal; (b) contacting the pregnant thiosulfateleach solution and/or a barren thiosulfate leach solution with a sulfideand/or bisulfide and/or a polysulfide to convert polythionates in thepregnant thiosulfate leach solution and/or barren thiosulfate leachsolution into thiosulfate; and (c) thereafter contacting the pregnantthiosulfate leach solution and/or barren thiosulfate leach solution withan oxidant to solubilize precipitated precious metal precipitates. 81.The process of claim 80, further comprising: recovering the solubilizedprecious metal from the pregnant leach solution.
 82. The process ofclaim 80, wherein in contacting step (c) the oxidant is molecularoxygen, the concentration of dissolved molecular oxygen in the pregnantleach solution and/or barren thiosulfate leach solution is at leastabout 1 ppm, and the pregnant leach solution and/or barren thiosulfateleach solution has a pH of from about pH 5.5 to about pH
 12. 83. Theprocess of claim 82, wherein in contacting step (b) the concentration ofthe oxidant in the pregnant leach solution and/or barren thiosulfateleach solution is no more than about 1 ppm.
 84. A process for recoveringa precious metal from a precious metal-containing material comprising:contacting the precious metal-containing material with a thiosulfatelixiviant at superatmospheric pressure to dissolve the precious metaland form a pregnant thiosulfate 5 leach solution containing thedissolved precious metal, wherein the concentration of added sulfiteduring the contacting step is no more than about 0.01 M; and recoveringthe solubilized precious metal from the pregnant thiosulfate leachsolution.
 85. The process of claim 84, wherein in the contacting stepthe thiosulfate lixiviant has a pH of less than 9 and a temperature offrom about 40 to about 80° C., and the total ammonia concentration inthe thiosulfate lixiviant is less than 0.05M.
 86. The process of claim84, wherein in the contacting step the thiosulfate lixiviant containsless than 20 ppm of copper.
 87. The process of claim 84, wherein thethiosulfate lixiviant includes sulfite and a total concentration ofsulfite in the thiosulfate lixiviant is no more than about 0.02M.
 88. Aprocess for recovering a precious metal from a precious metal-containingmaterial, comprising: (a) contacting the precious metal-containingmaterial with a thiosulfate lixiviant to solubilize the precious metaland form a pregnant thiosulfate leach solution containing thesolubilized precious metal, wherein the pregnant thiosulfate leachsolution includes polythionates; (b) contacting the pregnant thiosulfateleach solution with a reductant to convert the polythionate intothiosulfate; and (c) thereafter recovering the solubilize precious metalfrom the pregnant thiosulfate leach solution.
 89. The process of claim88, further comprising before the contacting step (b): separating atleast most of the precious metal-containing material from at least mostof the pregnant thiosulfate leach solution.
 90. The process of claim 88,wherein the reductant is at least one of a polysulfide, a sulfide, and abisulfide.
 91. The process of claim 88, wherein the pregnant leachsolution in contacting step (b) has a pH ranging from about pH 10 toabout pH
 11. 92. The process of claim 88, wherein in contacting step(b), the pregnant leach solution has a dissolved molecular oxygencontent of no more than about 1 ppm.
 93. The process of claim 88,wherein contacting step (b) includes the step of precipitating at leastmost of the solubilized precious metal to form a barren leach solutionand precious metal precipitates and the thereafter recovering stepincludes: separating at least most of the precious metal precipitatesfrom the barren leach solution; and thereafter recovering at least aportion of any remaining dissolved precious metal in the barren leachsolution.
 94. The process of claim 93, further comprising: (d) adjustinga pH of the barren leach solution to a pH of from about pH 7 to about pH12.
 95. A process for recovering a precious metal from a preciousmetal-containing material, comprising: leaching the precious metal fromthe material with a thiosulfate lixiviant to form a pregnant leachsolution including at least most of the precious metal in the materialand a metal impurity; recovering the precious metal from the pregnantleach solution to form a barren leach solution; and contacting at leastone of the pregnant leach solution and the barren leach solution with areductant to reduce a concentration of the metal impurity, therebyinhibiting a reaction between the thiosulfate and the metal impurity.96. The method of claim 95, wherein the reductant is one or more of asulfide other than a polysulfide, and a bisulfide, and a polysulfide.